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By John B. Patton

No academic program of reasonable duration can provide a geologist with all the skills that may be needed for applied work in industrial minerals, but any curriculum that does not provide background in such basic subdisciplines as mineralogy-petrology, stratigraphy, sedimentology, structure, and geomorphology will leave the practitioner handicapped. Enough geochemistry to understand materials requirements, and sufficient geophysics to utilize geophysical data, are essential. In addition to purely scientific considerations, the industrial minerals geologist must be able to evaluate economic factors and may be faced with legal, social, and political problems in land use, zoning, environmental protection, and reclamation.

Jan 1, 1981

By SME SME

List of 43rd ICGCM 2024 ORGANIZING COMMITTEE

Jun 25, 2024

By Roussos Dimitrakopoulos, Liam Findlay

Semimobile in-pit crushing and conveying (IPCC) systems can help reduce truck haulage in openpit mines by bringing the crusher closer to the excavation areas. Optimizing a production schedule with semimobile IPCC requires integrating extraction sequence, destination policy, crusher relocation, conveyor layout and truck fleet investment decisions. A mining complex with multiple mines and IPCC systems should be optimized simultaneously to find an optimal schedule for the entire value chain. An integrated stochastic optimization framework is proposed to produce long-term production schedules for mining complexes using multiple semimobile IPCC systems. The optimization model has flexibility to select the crusher locations and conveyor routes from anywhere inside the pits. The framework uses simulated orebody realizations to consider multielement grade uncertainty and manage associated risk. A hybrid metaheuristic solution approach based on simulated annealing and evolutionary algorithms is implemented. The method is demonstrated using an iron ore mining complex

Nov 1, 2024

By Q. Feng

Study has been carried out to apply fly ash as a high strength, water-resistant precast construction material through geopolymerization. Experiment results show that the working conditions such as water content, the concentration of NaOH, curing temperature and curing time, significantly affect the mechanical property of geopolymer matrix. Through optimization, an above 100 MPa compressive strength has been achieved with the geopolymerization products. Water soaking tests show that the geopolymerization product has a very high water resistance without losing noticeable compressive strength even after a one month soaking time. To elucidate the geopolymerization mechanism, microscopic techniques such as SEM/EDS and XRD are also applied to investigate the microstructure, the elemental and phase composition of geopolymerization products. The findings of the present work provide a novel method to apply fly ash as a high strength water-resistant precast construction material. INTRODUCTION Geopolymerization is a process of obtaining a polymeric structure from aluminosilicate by dissolving aluminosilicate sources in a strong alkali (NaOH) solution and condensation of free [SiO4] and [AlO4] tetrahedral units (Davidovits, 1985 and 1988; Duxson et al, 2005; Xu et al, 2000, 2002). After being cured at an elevated temperature for a specific time, the geopolymerization product shows a greatly increased compressive strength. Therefore, the product can be used as construction material such as bricks and road pavement. It provides a very promising method to apply aluminosilicate sources such as fly ash as a construction material for a sustainable development. Fly ash is a main industrial waste which needs to be carefully impounded because of its potential impact on environment. Many efforts have been tried to utilize fly ash as construction materials substituting Portland cement because of the following reasons. First, fly ash is a silica/alumina rich aluminosilicate source, which makes the specific material an ideal source for the geopolymerization process. Secondly, the particle size of fly ash is usually very small (< 50 microns) and it saves a lot of energy consumed during the production process of cement for grinding. Thirdly, fly ash is an ideal source for geopolymerization because it is reactive to alkali even at a mild temperature. In the past twenty years, many studies have been carried out on the geopolymerization of fly ash to a make construction material. However, there are some disadvantages associated with the conventional methods of treating fly ash (Palomo et al, 1999; Perera et al, 2007; Temuujin, et al, 2009; Hanjitsuwan et al, 2014). First, in most of the proposed geopolymerization methods, a salient amount of sodium silicate has to be added. It therefore makes it complex whether the achieved mechanical strength of geopolymerization products is due to a real geopolymerization process between fly ash and alkali or just largely the solidification of metal silicate in air at an elevated temperature. Secondly, the authors have noticed that when the geopolymerization product is soaked in water, the solution pH increases sharply and the products will lose its compressive strength gradually in water. Thirdly, the compressive strength of the geopolymerization product as reported is usually less than 50 MPa.

Jan 1, 2019

By CK Tsang, Jack Muir, Dharman Gersch, Joshua Barry

Correct face pressure is critical to safe operation of a closed face TBM. If inadequate face pressure is applied, this will lead to excessive ground movement, and may result in collapse of the tunnel face. Face pressure calculations are performed manually every few meters along the tunnel alignment. With the application of a 3D ground model, ground profile information can be extracted for automatic computation. This will significantly improve the efficiency and accuracy and is adaptable to changes in design parameters. This paper will discuss the development of automatic computation techniques in extracting information from a 3D ground model.

Jun 13, 2021

By T. J. De Mull

The Alligator Ridge Mine is located 70 miles northwest of Ely in White Pine County, Nevada. Development of the mine began in the late 1970&apos;s as a joint venture between Amselco Minerals, Inc. and Occidental Minerals Corporation, with Amselco as the operating partner. Early in 1983 Nerco Minerals Company acquired Occidental&apos;s share of the property. The mine and ore processing facility have operated at capacity since 1981 producing 680,000 tonnes of ore and 1,900 kg of gold per year. The gold ore deposits at Alligator Ridge consist of disseminated micron-size gold particles in a silicified siltstone matrix. The deposit comprises three separate but adjacent areas of mineralization that are designated as Vantage 1, Vantage 2, and Vantage 3 after the name of the basin in which the deposits lie.

Jan 1, 1983

By J. Rostami, P. P. Nelson, A. Khetwaln

Tunnel boring machines (TBMs) can be considered as the method of choice for tunneling in many projects due to the high level of efficiency and high potential productivity, safety, and applicability in a variety of subsurface conditions. The muck transportation system in a TBM tunneling operation plays a significant role in the process and impacts the overall performance of the machine. This paper introduces a recently developed method of simulating TBM operation through discrete event simulation (DES), which allows for tracking of the various activities, incurred downtimes in the operations, and overall machine utilization. The study presented here compares two different muck transportation systems, conveyor and rail, under similar geological conditions to quantify the impacts of the choice of muck transportation systems on machine utilization. The results show the applicability and extent of redundancy in the use of the two systems relative to machine capabilities and their sensitivity to geological conditions.

Jun 13, 2021

By U. K. Custred

In 1973, the United States production of marketable phosphate rock set a record in excess of 42 million tons. This production rate is expected to continue to increase, due to the growing international requirement for fertilizer, at a rate exceeding 5% per year well into 1977. This trend contrasts sharply with the reduced phosphate rock production levels of the 1966 - 1971 period. The natural consequence of this growing phosphate rock market is an evaluation of the variety of means whereby plant capacity can be most effectively increased. One approach toward increased plant production capacity is through large capacity flotation cells provided they can be successfully applied to the recovery of the relatively coarse product which is characteristic of the phosphate rock feed. This paper will summarize the results of a nine month evaluation test program, conducted at the Haynesworth Mine at Bradley, Florida in which the flotation circuit "specific capacity" (tons feed/ft.2 floor space) and "specific input power" (Hp/ton feed) were determined, and the feasibility of achieving economically acceptable concentrate grade and recovery levels was demonstrated, using large, high capacity, flotation cells.

Jan 1, 1974

By S. S. Peng

In this paper, the precise definitions for a number of commonly used subsidence parameters are given. The subsidence parameters have been determined using an integrated approach for 110 subsidence cases collected in the major US coal fields. Attempts have been made to establish empirical formulae for the parameters. The results of this study indicate that some traditionally-held concepts should be changed.

Jan 1, 1995

By C. F. Giermak

Over a considerable period of time great interest has been shown in selective dry concentration of fine ferromagnetic particles, although it is only recently that new permanent magnetic materials have been made available to permit building of such permanent magnetic separators. It is well known that the conventional drum separator does&apos; not efficiently separate fine materials. The centrifugal force on any drum acts in accordance with the particle size and gravity, which may beat varying magnitude-.opposite to the magnetic force, especially when the material is not closely sized. The Magnetic-Mechanical Separator (MMS) is designed for the concentration of fine, unsized magnetic materials where the percentage composition of magnetics to nonmagnetics, or the nature and texture of the materials, is such that its unique agitation and shake-out action with no centrifugal force provides greater efficiency than is obtainable by conventional magnetic separation equipment. This machine consists of a complete double belt conveyor unit with a premagnetizing Axial Pole Agitating Pulley at the feeding end of the lower conveyor belt and a special agitating concentrator magnet above the lower belt of the upper conveyor.

Jan 1, 1965

By S. Afrouz, C. Keles, E. Sarver, A. Greth

While respirable dust monitoring has long been done on the basis of mass concentration, it is increasingly recognized that specific dust characteristics may play an important role in health outcomes. Scanning electron microscopy with energydispersive X-ray (SEM-EDX) can be used for detailed particle analysis, but the quality of results can be dependent on the sample substrate; ideally, particles should be deposited on a smooth, uniform substrate like polycarbonate (PC). While direct air sampling onto PC filters is possible, this is not the standard approach in many situations. For example, in coal mines, polyvinyl chloride (PVC) filters are often used. They enable conventional mass-based analysis — of the total respirable dust, and the content of crystalline silica in this fraction — but are fibrous and thus not ideal for SEM-EDX or other microscopy. To address this issue, the current work tested a simple method to recover respirable dust particles from PVC filters and redeposit them to PC. Results generally indicated that recovered particles are representative of the original dust sample in terms of mineralogy distributions. Where this was not the case, the issue appeared to be with high particle loading density during SEM-EDX work rather than related to the recovery/redeposition method itself. This bodes well for applications where a single PVC filter sample could be used for nondestructive analysis to determine respirable dust mass and/or respirable silica mass, followed by preparation for SEM-EDX analysis to evaluate particle characteristics. The recovery and redeposition method might also be adapted for other applications, such as to enable the analysis of respirable dust samples collected by the continuous personal dust monitor (CPDM).

Oct 1, 2024

By L. C. Lang

INTRODUCTION The introduction of 165-mm (61/z-in.) holes to underground mining operations has made possible the application of Canadian Industries Ltd&apos;s (CIL) vertical crater retreat (VCR) mining method, illustrated in the accompanying sketches. This unique and revolutionary new application of spherical charge technology (see the Appendix), when applied to primary stopes and pillar recovery, eliminates raise boring, slot cutting, and dilu¬tion of ore by backfill; greatly improves fragmentation; reduces labor and time requirements; eliminates uphole drilling and blasting; and minimizes or completely elimi¬nates damages by blasts to the walls and retreating backs of the stope or pillar. If vertical large diameter holes are drilled on a designed pattern from a cut over a stope or pillar to bottom in the back of the undercut, and spherical charges of explosives are placed within these holes at the calculated optimum distance from the back of the undercut and detonated, a vertical thickness of ore will be blasted downwards into the previously mined area. Repeating this loading and blasting procedure, min¬ing of the stope or pillar retreats in the form of horizontal slices in a vertical upwards direction until the top sill is blasted and the mining of a stope or pillar is completed. The VCR method is also applicable to drop raises and has the potential to replace all other raising methods under most circumstances. PILLAR RECOVERY Levack Mine Inco Metals Co., Ontario Div., provided the first opportunity for the method in pillar recovery. The Levack mine&apos;s high grade ore pillar No. 4800 on the 975-m (3200-ft) level was used for the production¬scale experiment (Figs. 1 to 3). The pillar was about 49 m (160 ft) long, 6 m (20 ft) wide, and 20 to 26 m (65 to 85 ft) high. The mined area on both sides of the pillar was backfilled with a 1:30 cement:sand mixture. The pillar was removed in two phases. In phase 1, the standard uphole method was used to blast down the 18-m (60-ft) long section of waste from the bottom of the ore into the undercut. From the pillar&apos;s top sill, 165-mm (61/a-in.) holes were drilled downward into the pillar, and by measuring the depth of the holes, the results of the uphole blast were determined and roof line 1 was established. The bottom of each hole was plugged, then filled with sand to place the center of gravity of each spherical charge (loaded from the top sill) at a predetermined optimum distance from the horizontal free surface. The charges were then detonated. After detonation, both draw drifts were full of extremely well-broken material. The depth of each hole was measured again, and the plot of these depths re¬sulted in roof line 2. The same blasting procedure was repeated and the resultant new back elevation was marked by roof line 3. The poor results of the initial uphole blast at one location (notice the peak in area 1) appeared to influ¬ence the subsequent new backs. A third blast success¬fully evened the back, and resulted in roof line 4. An unblasted slab averaging 6.3 m (20.9 ft) thick remained below the pillar&apos;s top elevation as the final sill. In all three spherical charge blasts fragmentation of the blasted material was extremely fine. The backfill was fully exposed on both sides of the now-blasted pillar. The backfill remained undamaged and the ore was not diluted by sand. The remnants of all the 165-mm (61/2 -in.) holes remained clean and undamaged, and the holes had well-defined bottoms that could be easily measured and plugged. Each blast took down a 3.9 to 4.2-m (13 to 14-ft) thick horizontal slab of ore. Productivity was three times greater than that of the previously practiced cut-and-fill method. Since this was the first such experiment, blasting the remaining 6-m (20-ft) thick final slab was the sub¬ject of some deliberation. If the described method was repeated, we could have ended up with a 1.8-m (6-ft) thick sill unsafe to work on. It was therefore decided to blast the whole sill using two spherical charges prop¬erly placed in each hole, but with the application of

Jan 1, 1982

By R. L. Grayson, R. C. Althouse, L. A. Layne, M. J. Klishis

Introduction The researchers combine a detailed microanalysis of accident data and roof bolting field observations to develop new and meaningful risk indices for the roof bolter job classification. These indices incorporate information that has not been used in traditional measures of risk. The significant parameters that these new indices incorporate are the total number of accidents occurring within a particular work activity, the amount of time a miner usually spends performing the work activities in which the accidents occurred, and the average severity of the injuries incurred in the accidents. Being specific for a particular job, they aid comparisons among work tasks performed in the bolting cycle. The indices&apos; specificity provide anew basis for reexamination of training and operational development of miners for specific jobs. Microanalysis The roof bolting microanalysis was performed by examining the empirical data and written summaries for all bolting accidents reported in the West Virginia Safety Information System (WVSIS) for the years 1983-1987. The WVSIS database contained more than 15,000 accident narratives. The microanalysis of bolting accidents covered 2083 reported personal injuries (RPIs). This large number of injuries, performed at an industry-wide level, provides a broad picture of the nature of roof bolting injuries. The analysis focuses on the activities associated with the cyclical routines of the roof bolting process. For this analysis, roof bolting tasks were categorized into work routines based on the sequence of work activities performed by roof bolters. Four work routines were identified within the overall bolting cycle. And these routines comprised 70% (1217) of all roof bolting accidents at the face. The routines include: face area preparation; tramming, positioning and setting the automated temporary roof support (ATRS); drilling holes; and installing bolts. The types of injuries most common within a routine and the average severity of those RPIs were also identified. Table 1 gives a breakdown of the percentage of accidents by the bolting routines. Another 19% of the roof bolting RPIs cannot be classified in a particular routine because of the vagueness of the narratives. However, it was clear that these accidents involved drilling holes or installing bolts. These results are consistent with previous research. It identified drilling and bolting, followed by tramming, as the most frequent sources of injuries (Helander, Krohn & Curtain, 1983). [Table 1-Accidents by bolting routines at the face Accidents Bolting routine Percent Number Face area preparation 18.2 222 Tram, position, ATRS 23.3 284 Drilling holes 34.2 416 Installing bolts 24.2 295 Total 1217] Roof bolter operators were involved in 82% (1706) of the 2083 accidents. The RPIs incurred by classified roof bolters occurred largely on working sections (95%) and at the face area on the working section (75%). Approximately 23% of the 1706 roof bolter RPIs occurred during tasks which were not part of the bolting cycle. These included belt utility work, providing supplies and section moves. Miners in other job classifications who were involved in roof bolting accidents accounted for the remaining 18% of the 2083 roof bolting RPIs. Among this group were section utility men and general laborers (41%), section machine operators (38%), and electricians and mechanics (11%). Another 9% were performing on-section supervisory functions. Face area preparation Face area preparation work led to a substantial number of injuries, although it accounted for the lowest number of injuries of the four routines. Several tasks were associated with the injuries incurred in this routine. Table 2 shows the percentage of injuries associated with face area preparation. [Table 2- Accidents occurring during face area preparation 26.6% Setting temporary and supplemental supports 24.8% Scaling roof and ribs 21.2% Installing ventilation devices 15.8% Handling supplies and materials 11.6% Miscellaneous tasks] The injuries commonly involved overexertion and &apos;struck by&apos; falling coal and rock. Handling materials often led to overexertion injuries. Nearly 66% of the scaling-related injuries and about 33% of the injuries incurred during the setting of temporary supports involved &apos;strikes&apos; by equipment, materials, top and rib. Tramming, positioning and setting the ATRS During tramming operations, workers were often struck by components of the machine and hit by loose roof and rib. Poor footing also led to slips, sometimes against the moving machinery. Table 3 gives a breakdown of accidents in this routine. Control levers were a significant problem in this routine.

Jan 1, 1993

By R. T. Torrie, J. R. Mernagh, D. B. Chambers

l, INTRODUCTION Uranium mine and mill workers are exposed to external gamma and beta radiation fields from radioactive ore. In the past the average uranium content of ores mined in the United States and Canada ranged from about 0.1% to 1.0% U308 with pockets of much higher grade ore. Holiday (1973) reported that radiation surveys in the U.S. uranium mines found mean gamma "radiation rates ranged from 0.20 to 0.70 mrem/h. Such radiation rates cause relatively insignificant exposures." Others also concluded that the external gamma radiation fields associated with uranium mining did not result in significant worker exposures (Federal Radiation Council, 1967; Simpson, S.D. [et al], 1959) Gamma exposure levels in most modern-day Canadian uranium mines are reported to be low with average annual exposures estimated to be less than 1 rem/a (Frost, S.E. [et al], 1981). However, two developments are taking place which affect the potential significance of external radiation fields in the uranium mining environment. The first development stems from the most recent system of dose limitation developed by the ICRP which is intended to limit the workers&apos; overall risk from exposure to ionizing radiation through the adoption of a sum rule (ICRP 26, 1977) which combines external and internal radiation exposures. In the case of exposure to radon daughters the sum rule will have the effect of reducing the annual exposure limit below the recommended limit by an amount that depends on external radiation and other sources of internal exposure such as the inhalation of ore dust (ICRP 1980). The Atomic Energy Control Board of Canada (AECB) is reviewing this subject and is expected to produce its recommendations shortly. Irrespective of the form of the sum rule eventually adopted by the AECB, it is clear that the net effect of the sum rule will be a collective reduction of the individual dose limits for individual exposure pathways. The second factor is the increasing development of high grade uranium deposits in Northern Saskatchewan. Some of these ore bodies have an average ore grade of 1% to 5% U308 or greater. Since the potential external radiation fields increase in proportion to the ore grade, it is apparent that increased effort in radiation protection planning is required in order to develop safe yet workable methods for mining and milling such ores. This paper is intended to provide information which can be of assistance in the formulation and development of a mining and milling plan. The principal focus of the paper is source identification and the design of radiation protection measures to limit external gamma radiation exposure. The exposure of workers to external beta radiation fields is also discussed. The paper is organized as follows: - Section 2 deals with source characterization. - Section 3 discusses the effects of finite source size and distance (i.e. geometry effects). - Section 4 presents selected data that are useful in evaluating shielding requirements. - Section 5 discusses the potential beta radiation fields. - Section 6 discusses practical data requirements for worker exposure scenarios. - Section 7 presents a variety of work exposure calculations. - Section 8 is a summary of this paper. 2. SOURCE CHARACTERIZATION This section develops the basic formulae for estimating the fluxes and doses from external gamma radiation. The calculation of the radiation flux due to a distributed source (i.e. a linear, area or volume source) as a function of distance assumes that any distributed source can be treated as a summation of point sources. [ ] Uranium ore contains radionuclides from both the decay chains of U-238 and U-235. In this paper the radioactive daughters are assumed to be in secular equilibrium with the uranium parent. (If natural thorium were present in the ore in significant quantities, the gamma rays originating from Th-232 would have to be added to the gamma rays from the uranium series). In all, there are over 50 separate gamma rays (as well as alpha and beta particles) emitted from the U-238 and U-235 radioactive decay series (USHEW 1970). The total

Jan 1, 1981

By O. A. Bascur

In recent years, metal-producing companies have increased their investment in automation and technological innovation, embracing new opportunities to enable transformational change. Transformation to a digital plant can fundamentally revolutionize how industrial complexes operate. The abundant and growing quantity of real-time data and events collected in the grinding and flotation circuits in a mineral processing plant can be used to solve operational issues and optimize plant performance. A grade recovery model is used to identify the best operating conditions in real time. The strategy for increasing the value of instrumentation in current plants is reviewed. An optimal Gaudin size distribution model provides augmented information from traditional sensors to find the optimal grind cut size to reduce metal losses in the flotation circuits. Sensors in flotation circuits enable an estimate of the recovery and determination of the optimal froth depth and aeration using an air hold up flotation model. A strategy of classifying data for on-line generation of insights to using operational intelligence tools is described. The implementation of a recovery/grind strategy with industrial examples in non-ferrous mineral processing is presented.

By S. Bhattacharaya, S. Aich, B. K. Nandi

"At present scenario, there has been an increasing demand in the usage of blend coal in power sector; regardless of how, specific knowledge on burning and ignition characteristics and their interactive effects is stingy. In this work a detailed investigation was therefore taken up to study in detail the combustion characteristics of selected coal samples and their blends. Certain parameters were believed to have significant effects on combustion characteristics of individual coal and their blend. Major parameters identified for this purpose included proximate analysis, combustion conditions, heating rate, reaction regions, blending ratios, ignition condition, reactivity (or reaction rate) and burnout temperatures. With the help of these parameters, thermal features and reaction kinetics of the individual and blended samples were investigated through a qualitative and quantitative thermo-gravimetric analysis of two different individual coals and their blend carried out in an oxygen atmosphere from ambient to 800 0C at a constant heating rate of 10 K/min. A significant difference in terms of rate of mass loss with temperature correlated from Arrhenius expression along with coal-coal blend interactions and deviation from their expected behaviour, have been observed between the kinetics of individual and blend samples, the latter showing better ignition performance and comprehensive combustion phase. It was also found that particles having the lowest percentages of volatile matter and ash could possibly be blended with particles having high ash and volatile matter to meet the required burning condition which can be effective in terms of cost savings, improved fuel economy, reduced undesirable exhaust emissions and improved safety of coal burning equipment. Experimental results further indicated that individual samples and coal blend samples have different combustion features in terms of residue left, peak temperatures and activation energies. Thermodynamic activation parameters such as activation energy, pre-exponential factor and order of reaction were computed from the thermal data using the modified form of non-isothermal methods. INTRODUCTION Coal has been the most significant source of energy for India and would remain so at least for the next few decades. About 60-65% of the electricity produced in India is from coal fired power plants. Indian coal is of drift origin and of high ash content. Ash content of the coal burnt in power plants in India is in the range of 30-45%, most of it being Run-of-Mine coal. Most of the power plants in India source their coal from number of mines, most common being multiple sourcing. Usually coal from each source [1] is burnt separately. For various reasons mostly associated with logistics, burning of blended coal is rarely practiced. Blending of coal is practised mainly for improving the quality of available coal and meets the shortage of coal supply. Blending technique is generally practised in power industries to reduce the cost of power generation along with mitigating combustion performance or slagging problem ensuring consistency in as fired coal quality and also for environmental compliance. Blending good quality imported coal can improve coal quality and helps in preserving indigenous coal resources for longer period [2]. Use of blended coal shall reduce load on all boiler components (Mill, CHP, Boiler, AHP) & reduce O&M cost and increase the availability. The main purpose of blending is to get a blended as fired coal having ash, volatile matter and gross calorific value close to design coal and the two coals are to be blended in such ratio proportions that resultant coal blend have desired ash, volatile matter and gross calorific value [3]. For a coal to be blended certain parameters have to be considered to get compatibility with respect to combustibility ( Burn ability) of coal, grindability ( HGI), ash fusion behavior, swelling characteristics of coal, need to be checked"

Jan 1, 2017

By Ingemar Marklund

INTRODUCTION Today at LKAB, the primary mining method used is sublevel caving, independent of ore-body size, shape, dip, and geographical location. The mining of small ore bodies with relatively flat dip involves a lot of develop¬ment work, a large initial investment, big dilution, and large ore losses. In 1975 serious investigations into other possible methods, especially for those small ore bodies which do not lie in direct contact with other ore bodies, began. Different methods were studied and a number of in¬teresting ideas came from Russian mines. An analysis was made of one of the proposed methods, sublevel benching, in which mining is carried out from raises, and which uses the throw from the blast to place the broken ore. However, there were reservations about working in big open rooms at great depth, and doubts about the possibility of getting acceptable drilling results in an acceptable environment from raised platforms. Testing was therefore delayed for a time. However the investigations continued and in the fall of 1976, it was decided to try vein mining in a small ore body (Indian) (see Figs. 1 and 2). MINING RESULTS The results are now available from the mining of Indian and are as follows: Ore plan area 1200 m2 Level height 50 m Ore volume 60 000 m3 Ore tonnage 270 000 t Iron content 62% Dip 0.69 rad (40°) Ore thickness 14 m Ore compressive strength 26 MPa Ore C-factor (fracture density) 0.21 Hanging wall compressive strength 28 MPa Hanging wall C-factor 0.27 About 40 000 m of the holes were drilled by ring drills with an average productivity of 85 m/shift; loading of a round (2 rows, 750 m of hole) was usually accomplished in one shift. The total ore recovery was about 270 000 t with an average iron content of 59%, and three to four shifts were required per round (4700 t) to load. The principal of "throwing" the rock during blast¬ing worked well, and no ore remains on the footwall. ROCK CONTROL Before mining commenced at Indian, it was dis¬covered that the rock strength in its vicinity was rela¬tively poor. Since the ore is flatly dipping, and mining was to take place with open rooms, the strength of the hanging wall was of great importance. A supplementary program of exploration drilling was done during the fall of 1976 in order to assess fracture spacing and rock strength from the drill cores. Results showed that the strength of the hanging wall was low because of the many fractures. A decision was made to try cable bolting of the hanging wall by extending some of the ordinary blast¬holes. Cables (20 mm diam) were grouted in these larger holes and the rest of the hole was then utilized for ordinary loading (see Fig. 2). About 100 bolts altogether were installed and, to date, only small falls have occurred and the hanging wall is holding. Its exposed area is 2100 m2 DEVELOPMENT WORK On the 350-m level, four crosscuts (two per mining room), one longitudinal drift, and two elongated drifts (one per raise) for the raise climber were developed as shown in Fig. 1. Of the 230 m of drift, 60 m were in ore and the rest in waste. For mining room A alone, a total of 115 m of drift were needed with 30 m in ore. The drift dimensions were about 4.5 X 5.5 m. In the drifting on the 300-m level, two connecting drifts were driven from one existing transport drift forward to the respective raises. The total development was 93 m, all in waste (Fig. 1). The two raises having a length of 80 m each were driven at an angle of 0.68 rad (39°) from the hori¬zontal. They were driven along the footwall contact with a separation of 30 m and had an area of 10 m2. LONG-HOLE DRILLING The drilling of long holes with the ring drilling unit began in the fall of 1977 in raise A. The equipment consisted of an Alimak raise climber, equipped with a lower platform from which the charging was done.

Jan 1, 1982

By Janet A. Johnson, T. B. Borak, K. J. Schiager

INTRODUCTION In the USA regulatory agencies have adopted standards which limit the allowable concentration of ore dust in underground uranium mines. The American Conference of Governmental Industrial Hygienists has recommended a threshold limit value for silica which has been incorporated by reference in federal regulations (30 CFR 57, 1980). The Nuclear Regulatory Commission (NRC) imposes a limit for unprocessed uranium ore dust for reducing the exposure to radioactive materials from inhalation (10 CFR 20, 1980). We have been unable to trace the origin of the NRC standard. Recently the ICRP has published recommendations concerning limits for intakes of radionuclides by workers (ICRP, 1979). We have used the ICRP methodology to compute the dose equivalent commitment to lung and bone from inhalation of insoluble ore dust particles. These values were used to derive air concentrations which would provide an acceptable risk for induction of cancer from occupational exposure. Results indicate that the new ICRP recommendations are more restrictive than the present NRC standard for radioactivity. The maximum allowable dust load in a uranium mine will depend on both the concentration of silica (% Si02) and the amount of radioactivity (% U308). We have combined these standards assuming complete independence of biological hazard to establish boundary conditions for which either silica or radioactivity is the dominating factor. Based on these results we present suggestions for analysis of ore dust to insure compliance with regulatory agencies. SILICA STANDARDS Silicosis, fibrotic lung disease produced by crystalline silica, was one of the earliest forms of occupational disease to be recognized. The pulmonary lesions caused by silica, silicotic nodules, consist of concentrically arranged bundles of collagen fibers. Fusion of these nodules results in progressive massive fibrosis causing alveolar changes which decrease ventillation and blood flow in the lungs. The current Threshold Limit Value (TLV) for mineral dusts is based on the concentration of free silica. The standard is expressed in three forms: (1) Particle count standard: TLV (mppcf) = 300/(% Si02 + 10) (mppcf = million particles per cubic foot) (2) Respirable dust mass standard: TLV (mg/m3 ) = 10/(% respirable Si02 + 2) (3) Total dust standard: TLV (mg/m3 ) = 30/(% Si02 + 3) The most commonly used form is the respirable mass standard. These curent values for silica are based primarily on epidemilogical studies of Vermont granite workers and other occupationally exposed workers. In the Vermont study no cases of silicosis were seen in individuals employed after dust control measures were initiated and whose subsequent exposure averaged less than 5 mppcf. The free silica concentration in the airborne dust to which these workers were exposed averaged 25%. A concentration of 10 mppcf of granite dust is considered equivalent to 0.1 mg/m3 quartz. The current respirable mass standard allows a maximum silica concentration of 0.1 mg/m3. NIOSH (1974) has recommended a reduction in the current TLV for respirable silica to 0.05 mg/m3 Si02. As with radiation standards, the TLV for mineral dusts is an upper limit for time weighted average exposure. It is recommended that dust concentrations be maintained as far below the TLV as current practices permit. RADIATION STANDARDS The present concentration limit for airborne natural uranium in ore dust prior to chemical separation is 75 µg of uranium per m3 of air ( 10 CFR 20, 1980) . This corresponds to 1.85 Bq/m3 (50 pCi/m3 ) of natural uranium, U-238, U-235 and U-234 and is equivalent to 3.7 Bq/m3 (100 pCi/m3 ) of gross alpha activity with radioactive equilibrium between the long-lived alpha progeny through Ra-226. The origin or basis for this standard is not known, but it presumably includes both chemical toxicity and risk from somatic radiation injury. Recently the International Commission on Radiological Protection (ICRP, 1977) has published recommendations concerning the objectives and criteria for limiting radiation exposures. This was followed b y a revision of the limits for intake of radionuclides by workers (ICRP, 1979; ICRP, 1980). The previous compilation of limits for internal emitters was published by ICRP Committee II in 1959 (ICRP, 1959). The latest version includes new methodlogy outlined in ICRP Publication 26 (ICRP, 1977) as well as extensive metabolic information

Jan 1, 1981

By Rafael Moure

On April 21, 1980 the Oil, Chemical and Atomic Workers Union (OCAW) submitted a petition urging the revision of the Mine Safety and Health Administration (MSHA) permissible exposure limit to radon daughters in underground mines. This petition asks for: 1. Annual worker exposure to be lowered from the current 4 WLM/year to 0.7 WLM/year; 2. Full economic protection for workers whose exposure exceeds the proposed 0.7 WLM/ year; and 3. The primary means of lowering workers&apos; radiation exposures shall be environmental controls. I will present the scientific rationale for the lower standard requested in our petition to the Department of Labor. [1. Radiation Exposures to Uranium Miners] Two types of radiation exposures are present in the environment of uranium mines: A. [External Exposure] to gamma sources which until 1978 was considered as negligible compared with internal alpha radiation. R. E. Utting (1), in an investigation of three Canadian uranium mines, found levels of exposures adding to an annual dose equivalent of 3 rems (30 mSv)/year. He concluded that gamma dosages may be a major fraction of the maximum permissible dose for many individual miners. The U.S. experience has shown that there is significant gamma exposure in U.S. uranium mines and, therefore, it could contribute to the total radiation exposure to the lungs in combination with internal alpha radiation. Before ventilation improvements in uranium mines, the contribution of external radiation to total lung dosage was relatively small. After improvements in ventilation, the reduction in internal radiation exposure appears to be in 1980 as low as 1/10 of the 1950 dosages. However, ventilation improvements do not affect external gamma radiation. Gamma radiation remains at the same 1950&apos;s level, and it represents today a higher proportion of the total radiation burden to the miner&apos;s lungs. MSHA regulations (57.5-47) state that if gamma exposure is below 2.0 milliroentgens per hour, then workers are not required to wear gamma dosimeters and cummulative gamma radiation records need not be kept. Assuming an exposure of 1.9 milliroentgens per hour per 173 hours/month for 12 months,a miner would accumulate a dose from external radiation of 3.9 rems/ year. This dosage would, of course, go unrecorded. MSHA, therefore, should consider a downward revision of the external radiation permissible exposure limit of 2.0 milliroentgens/hour. MSHA should make clear that this exposure level, when added to the internal exposure, does not make for sound health physics -even though the measured current levels do not approach 2.0 milliroentgens/hour. B. [Internal Exposure] to alpha radiation is still the larger contributor to lung irradiation in uranium miners. The target organ is still the lung. The parameters to be considered in determining the biological radiation load appear in Table I for radon daughters present in the upper respiratory tract (URT). TABLE I RADIATION DOSE PARAMETERS IN THE URT OF URANIUM MINERS Parameter Uranium Underground Mining [1. Quantity of the Absorbed Radon daughters (Po-218, Nuclide Pb-214, Bi-214) deposit in bronchial and lung tissue. (Measured by air concentration, Working Levels) 2. Type of Radiation Emitted Alpha Particles 3. Energy of Radiation High-Linear Energy Transfer 4. Half-life of the Nuclide Ra 222, 3.8 days in the Organ Po-218 (RaA), T1/2 = 3.1 minutes Pb-214 (RaB), T1/2 = 26.8 minutes Bi-264 (RaC), Vi = 19.7 minutes 5. Selection Factor of the Direct inhalation with mine Organ dust to be deposited in the tipper Respiratory Tract 6. Mass and Form of the Organ Especially cancer sensitive, tracheobronchial region and bronchi* 7. Distribution of the Activity Dependent on deposited dust in the Organ distribution in the Upper Respiratory Tract 8. Chemical Nature of All compounds (RaA, RaB, RaC) Incorporated Substance biologically insoluble. (Only localized, irradiation) *See Reference (2)] The parameters of Table I depend, in turn, on the demography of the exposed population, especially: 1. Age 2. Sex 3. Health Status 4. Genetic Constitution 5. Lifestyle (Smoking, Diet) Some of these demographic characteristics of uran-

Jan 1, 1981

By William D. Cassity, Batric Pesic

Introduction Gold mining on an industrial scale is becoming increasingly dependent on low-grade refractory sulfide ores that contain large amounts of pyrite (FeS2) and arsenopyrite (FeAsS). Gold is usually present as sub-microscopic particles dispersed evenly throughout the sulfide matrix. Without some form of preoxidation, efficient recovery of this finely disseminated gold through cyanide leaching is hampered because the surface of the sulfide particle becomes passivated and impervious to penetration of the cyanide. Traditional methods for recovery of refractory gold from arsenopyrite ores include roasting and autoclave leaching. A modem alternative method is to preoxidize the mineral using the bacterium Thiobacillus ferrooxidans prior to attempting to re- cover the gold through cyanidation. The bioleaching of arsenopyrite ores presents a problem in the mobilization of large quantities of arsenic, present both in the +5 and +3 ionic states. Dissolved arsenic species pose a dual problem: they are -1 to the bioleaching process itself through inhibition of bacterial activity, and they also pose an environmental threat. The arsenate anion (AsO2) has been shown to be more toxic to living organisms, including Thiobacillus ferrooxidans, than the arsenate anion (AsO43) (Collinet and Morin, 1989). There is some disagreement as to the fate of arsenite in bioleaching solution. Mandl, Matulova, and Docekalova (1992) reported that dissolved arsenite concentration was constant during a long-term bioleaching study of chalcopyrite using Thiobacillus ferrooxidans. Torma and Oolman (1992) report that Thiobacillus ferrooxidans has been found capable of oxidizing As3+ to As5+. The solid products of arsenopyrite/pyrite bioleaching include jarosite (MFe3(SO4)2(OH)6 where M can be K+, Na+, NH4+ or H3O+), scorodite (FeAsO4.2H2O), ferric hydroxides, and ferric hydroxysufites (Carlson et. al., 1992; Van Breemen, 1982; Lazaroff et. al., 1982). The particular species formed is a function mainly of pH. Little study has been performed on the effect of solid products on bioleaching. Pesic and Kim (1993) showed that Thiobacillus ferrooridans cells served as a nucleation sites for jarosite particles, which grew rapidly and eventually killed the cell. Most bioleaching studies are performed at low pH&apos;s (<2.0) under uncontrolled conditions. By studying bioleaching of arsenopyrite under controlled pH conditions and at higher pH&apos;s (2.0 to 3.0), the distribution of iron and arsenic between the solid and liquid phases can be controlled, and their effects studied. The objective is to develop a bioleaching process that operates at high efficiency while keeping amounts of dissolved heavy metals low. Pesic, Stohok, and Torma (1993) demonstrated the usefulness of this approach in the leaching of cobaltite (CoAsS) concentrates. Materials and Methods The focus of this study was the bioleaching of arsenopyrite ore by Thiobacillus ferrooxidans. The parameters studied included the effects of pH, and the effects of arsenate (AsO43 and arsenite (AsO2). pH studies included pH set initially but not controlled, and pH controlled continuously with NaOH or LiOH. Bioleaching results from two strains of Thiobacillus ferrooxidans with different adaptation histories were compared. Bioleaching experiments were conducted in 150 ml stirred glass reactors using a strain of Thiobacillus ferrooxidans adapted to arsenopyrite ore, supplied by the U.S. Bureau of Mines. These bioreactors were innoculated with stock cells grown on arsenopyrite ore for 14-16 days. Typical elemental breakdown for this ore was: As, 9.90%; Fe, 18.60%, Stot, 9.26%; Suifide, 7.50%; SiO2, 25.90%. Mineralogically, the ore contained both arsenopyrite and pyrite. Bioleaching efficiency was measured by withdrawing bioleach solution samples at selected intervals, filtering, and measuring the amount of dissolved cobalt, iron, and arsenic by atomic absorption spectroscopy. Because the precipitation of iron and arsenic at higher pH&apos;s would lead to unreliable results, a 100 mg tracer of cobaltite concentrate from the Blackbird mine in southern Idaho was added to solution. Dissolution of the cobaltite tracer was used to indicate bioleaching of the bulk arsenopyrite ore. Cobalt has been shown to be stable in solution even when iron and arsenic were precipitated (Pesic, Storhok, and Tonna, 1993). The pa-

Jan 1, 1995